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ABSTRACT T This paper identifies physical conditions responsible for past instability of the Main Hill area in Kennecott Copper Corporation's Bingham Canyon Mine, Bingham Canyon, Utah, and describes successful stabilization techniques employed in this slide zone. Geotechnical highlights of recent slides are presented, including monitoring, geological engineering data collection, hvdrology, physical rock properties, and stability analyses. Local conditions responsible for slope movement in the Main Hill zone are: (1) high pore water pressure that causes a substantial decrease in real and effective material strengths; 2) continuous structural systems that allow sliding in the more competent rock zones; and (3) broad areas containing fractured, altered, and incompetent rock and soil-like materials. The failure mode is structurally controlled and involves a circular-type soil failure in the toe region initiated by excessive pore water pressure. A surface mapping method is presented which assigns rock mass quality on the basis of block size and rock substance strength. This site characterization technique has successfully- predicted slide boundaries and failure modes; used in conjunction with laboratory determined shear strengths, it has provided basic input for calculating slope safety factors in the Main Hill slide zone. N The Main Hill slide zone is located in the northwest corner of Kennecott Copper Corporation's Bingham Canyon mine, Bingham Canyon, Utah. This particular area has been the site of several slope failures in past years. Significant movement occurred in 1956, 1967-68, 1974-75, and one to three bench failures have taken place repeatedly near the 6340 bench elevation. Remarkable similarities have been noted between the various large slides as to toe elevation, path of movement, major structural control, rock qualities, groundwater conditions, and slide width. Geological and engineering investigations of the Main Hill area have been done in-house and by consultants on several occasions. This paper presents geotechnical highlights of recent failures, including slide monitoring, geological engineering data collection, hydrology, physical rock properties, stability analyses, and stabilization techniques employed in the Main Hill slide zone. S Major movement in the Main Hill area occurred in 1967 and again in 1968. Indications of instability were first observed during Spring 1964. By March 1966, the slide boundaries were well established with tension fractures near the top of the slide zone at the 6820 foot elevation and basal thrusting at the 6340 level. Movement continued until Spring 1967, at which time outward displacement of the pit wail was very apparent as shown in Fig. 1. Accelerated movement began on June 18, 1967 and terminated late the following day. This activity displaced approximately two million tons of material between the 6900 and 6140 levels. The width of the slide ranged from 200 to 500 feet. The slide consisted of two parts: a relatively intact upper portion extending from a headwall crack on the 6900 level to a toe area on the 6620 level, and a rubblized lower portion extending from an intermediate headwall crack on the 6660 level to the toe on the 6140 level. A zone of slumping, containing some 300,000+ tons of highly fractured material, extended above 6900 to the 7020 level but was not directly involved in the slide movement.
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Geology > Rock Type (0.95)
- Geology > Mineral > Native Element Mineral > Copper (0.83)
- Well Completion > Hydraulic Fracturing (1.00)
- Reservoir Description and Dynamics > Reservoir Characterization > Reservoir geomechanics (0.87)
- Well Drilling > Wellbore Design > Rock properties (0.54)
ABSTRACT N Mining activity and basic geometry of open pit mines represent major obstacles in geotechnical mapping. Mining creates a constantly changing topographic base which may require corresponding updates of field sheets before mapping can proceed. The rapid pace of mining operations may also threaten removal of important geologic and engineering features before adequate base maps and surveyed control can be established. Maintaining such control during the mapping process is yet another complication. In addition, basic geometry yields nearly vertical outcrops. Access required for direct observation and measurement, essential to conventional plane table or Brunton techniques, may be impossible on such surfaces. Also, nearly vertical surfaces appear foreshortened in plan view. This reduction in area provides little room for recording geotechnical detail. The difficulties encountered in open pit mapping can be substantially reduced by utilizing a modified form of plane table photogrammetry. This technique employs photographs taken from the ground as the base upon which details are recorded. The information is later transferred to plan maps using the plane table principal of location by intersection (Davis and Foote, 1953). Any change in topography can be accommodated by simply taking additional photographs. Proper spatial representation in plan view can be determined later when updated maps and surveyed control are available. In addition, horizontal photographs from the ground portray vertical surfaces as they are normally viewed thus providing maximum area for recording details. Any detail which can be identified on the photograph can be located correctly even though the actual feature is physically inaccessible. The technique presented in this paper was developed by the author for special applications at Kennecott's Bingham operation in 1970. The procedure was rapidly adopted for general use and has proven to be a practicaltool for detailed investigation of potentially unstable areas (Zavodni and McCarter, 1976) and routine geological mapping (Richardson, 1975). Major advantages, other than those mentioned above, include: The speed and accuracy are greater than for conventional plane table or Brunton methods. The photograph represents a view from a station at some distance from the area being mapped. This vantage point allows evaluation of possible continuity of structures from one level to another which may not be evident at close range. The photographic base provides a permanent record of the topography at the time of mapping which permits re-evaluation of interpretations subsequent to mining. The technique does not require sophisticated and expensive equipment and can be applied with very limited knowledge of photogrammetry. S Before proceeding with a description of the technique, a few basic principles should be reviewed. Figure 1 represents the geometrical relationships between an object in the field and its image as photographed from the ground. The figure shows a typical horizontal photograph and a plan view of the camera and mining bench. The control point at A is brought into focus and recorded on the negative plane at point A". The line OA represents the distance from the camera station at point 0 to the field position of point A, and the line C"O represents the focal length of the camera.
- Materials > Metals & Mining (1.00)
- Energy > Oil & Gas > Upstream (1.00)
- Reservoir Description and Dynamics > Reservoir Characterization (1.00)
- Data Science & Engineering Analytics > Information Management and Systems > Artificial intelligence (0.54)
ABSTRACT ABSTRACT This paper discusses general principles of underground opening design based on the concept that with the necessary input data regarding the geology of a sine, the physical properties of the rocks, the in situ stress field in the rock and the general geometry of the underground openings, the rock formation can be classified into six major categories which are useful for design purposes. Using simple theories for the criteria of failure for rock in compression and tension, general elastic theory for stress concentration around openings and. in structures, beam theory for stability of roof rock, and elementary inelastic theory for creep phenomena, basic design techniques are outlined for the various rock classifications. The need for field evaluation of the theoretical design is stressed and some general techniques that should be used in these field evaluations are discussed. Where appropriate, the need for additional studies and research is noted. INTRODUCTION The basic theme of this session is on underground opening design; therefore, it seems appropriate that one paper should be devoted to a review of some of the general design techniques that have been developed and are in current use. Because of my long association with the Bureau of Mines, this paper is strongly influenced by the research efforts of my many colleagues. The general outline of the paper is based around lecture notes for a course in rock mechanics that I taught for several years at the Colorado School of Mines. This paper outlines some general principles and procedures that a rock mechanics engineer can use to arrive at a logical design for underground openings in various types of competent rock formations. The concept is used that rock formations can be classified into six major categories which are useful for design purposes provided the necessary, input data are available regarding the geology of the site, the physical properties of the rocks, the in situ stress field in the rock and the general geometry of the underground openings. Using sample theories for the criteria of failure for rock in compression and tension, general elastic theory for stress concentration around openings and elementary inelastic theory for creep phenomena, basic design techniques are outlined for the various rock classifications. Because even competent rock is not a perfectly elastic, homogeneous and isotropic medium, it should be understood at the very beginning of this discussion that the stability and safety of any theoretically designed rock structure is only an estimate of the true stability and safety of the mined structure. Thus, it is essential that fairly large safety factors (2 to 8) be employed in the theoretical design and that field evaluation of the design be made by instrumented studies during the construction of the underground structures. Table 1 lists what I consider are the basic requirements for the design and evaluation of safe and stable structures in rock. The first two items in this table involve the general geology of the site and are usually supplied by a geologic report. If the report is based on exploration drilling, it should contain a log of the structural defects in the core and an RQD index for the rock (Obert and Duvall, 1967).
- Materials > Metals & Mining (1.00)
- Energy > Oil & Gas > Upstream (1.00)
- Well Drilling (1.00)
- Reservoir Description and Dynamics > Reservoir Characterization > Reservoir geomechanics (1.00)
ABSTRACT INTRODUCTION The Climax Mine of AMAX's Climax Molybdenum Division is currently producing approximately 34,000 tons of molybdenum ore per day from two underground levels (see Figure 1), and initial design is underway for the next deeper (1200) level. The aim of this report is to present the methods of site evaluation used at Climax for cavability determinations and underground support design. Climax succeeds with its panel caving system of mining (Julin, 1973)(Figure 2) largely because of the presence of pervasive geologic structural discontinuities. The stock work ore body at Climax was formed by the multiple intrusion of mid-Tertiary rhyolitic magmas into Precambrian granite schist and gneiss and the hydrothermal emplacement of molvbdenite into the network of fractures that were generated during the forceful upward intrusion of magma. Subsequently, intrusive events that were bore deeply seated refractured the overlying, orebearing stock and propagated a complex system of axial and radial fractures. Post-ore movement along the Mosquito Fault (Wallace, 1968) and its associated stranded shear zones further fractured the ore body and disordered the existent geologic structure. The result of these structural events is a brittle, crystalline ore body that is transected by a complex pattern of faults, shear zones, and joints that augment cave initiation and enhance communication of the ore blocks. Generalizations can be made about the structural trends within the ore deposit (King, 1946; Kendorski, 1973), however, the intensity and orientation of geologic structure is by no means uniform throughout the pre body. The relative intensities of silicic and argillic alteration are also highly variable (Vanderwilt and King, 1955). As a consequence, no single rock-mass model suffices as a standard for mine design and site characterization for cavability and support design is a continuous process. SITE CHARACTERIZATION: A rock mass that caves readily is also a mass that is generally not conducive to the formation of stable underground openings. The two greatest concerns in the rock mechanics program at Climax are for the stability of the semi-permanent production drifts on the two existing working levels, and for the general cavability and stability conditions that can be expected for the future, deeper production levels. Site characterization is the major site requisite in approaching both problems: the main differences between site evaluation for the present and future production areas are ones of scale and detail. For the deeper levels, we are seeking to define the general cavability and stability of different sections of the mine. In our investigations on this larger scale, cavability and stability are considered to be inversely related, i.e., ground that is the most readily cavable is ground that is generally less structurally sound and more prone to develop stability problems, whereas ground that is more difficult to cave is ground that is also more structurally competent and less likely to need supplemental artificial support For initial future-level planning, the competence of the rock mass is taken to be the controlling factor in site characterization. For the present working levels, however, we are concerned with the stability of specific production openings, and the generalization of an inverse relationship between cavability and stability based solely upon a rock-mass competence rating cannot be directly applied.
- Geology > Rock Type (0.88)
- Geology > Geological Subdiscipline > Geomechanics (0.36)
- Materials > Metals & Mining > Molybdenum (0.88)
- Energy > Oil & Gas > Upstream (0.66)
- Production and Well Operations (1.00)
- Reservoir Description and Dynamics > Reservoir Characterization (0.89)
ABSTRACT INTRODUCTION This paper describes part of a geotechnical study based on stress determinations and convergence measurements conducted to assess the long-term stability of an experimental room-and-pillar oil shale mine. A rock mechanics program was initiated in 1971 at the Colony Mine as part of a large study conducted for the design of a commercial operation. The Colony Development Operation is a joint venture currently consisting of four active members; Ashland Oil Inc., Atlantic Richfield Company ( Operator of the project), Shell Oil Company, and The Oil Shale Corporation. Although oil shale commercialization plans are presently suspended, rock mechanics instrumentation has been continued as part of Colony's policy in maintaining readiness to proceed with a commercial shale oil plant when national energy policies became better defined. The property is located in the southern edge of the Piceance Creek Basin in northwestern Colorado, approximately 200 miles west of Denver. Experimental mining operations were conducted from 1965 to 1972 with one of the major objectives of the pilot mine being the assessment of opening and pillar sizes for the determination of extraction ratio and life of the property. Mining was conducted in a 60 foot thick portion of the Mahogany Zone in the Parachute Creek Member of the Green River Formation, at depths of 600 to 850 feet by a one-bench system. Pillar dimensions were 58 feet by 58 feet and rooms were 55 feet wide. A well defined system of joints and bedding planes is present in the flat lying oil shale beds. The pillars at Colony have two joint sets approximately 90 degrees to each other. The mean strike of the major set is E-W, with 42 percent of the dips being vertical ยฑ 10 degrees and 33 percent dipping south at 24 ยฑ 6 degrees from the vertical. The other dips range between these two orientations. Major pillar slabbing has taken place along the south-dipping joints. Joint spacing may vary from one foot to more than 15 feet. The minor joint set dips mostly vertically and has a wider spacing than the major set. The roof shows only one joint system parallel to the major pillar joint set. Most joints are small and very tight, but a small percentage are open and can be traced across some pillars. Joint filling material when present consists mostly of calcite. Laboratory tests indicate that the uniaxial compressive strength is dependent on oil content. Above 30 gallons per ton, strength seems to remain constant at 13,000 psi, but below this grade the strength increases with a decrease in oil content. The immediate roof rock is 3,000 psi stronger than the mine horizon because of a marked difference in grade. Field instrumentation was described in detail in another pape1r and basically consisted of in-situ stress determinations by the overcoring technique, and rock mass displacement measurements by means of borehole and tape extensometers. Design data was greatly enhanced by unplanned pillar and roof failures. PILLAR STRENGTH DETERMINATIONS Overcoring measurements were made before, during, and after pillar failures. These measurements indicated that average pillar strength was approximately 3100 psi.
- Geology > Rock Type > Sedimentary Rock > Clastic Rock > Mudrock > Shale (1.00)
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Geology > Rock Type > Sedimentary Rock > Organic-Rich Rock > Oil Shale (0.88)
- Reservoir Description and Dynamics > Unconventional and Complex Reservoirs > Shale oil (1.00)
- Reservoir Description and Dynamics > Reservoir Characterization > Reservoir geomechanics (1.00)
Fracture Propagation In Rock: Laboratory Tests And Finite-Element Analysis
Ingraffea, Anthony P. (Department of Civil, Environmental, and Architectural Engineering) | Heuze, Francois E. (Department of Civil, Environmental, and Architectural Engineering ) | Ko, Hon-Yim (Department of Civil, Environmental, and Architectural Engineering ) | Gerstle, Kurt (Department of Civil, Environmental, and Architectural Engineering )
ABSTRACT T Classical and new concepts of fracture mechanics are combined into a proposed method for the study of compressive fracture induced failure of rock structures. The strain energy release concept and a critical strain energy density fracture theory are applied to mixed mode fracture in simple structures. The theory is implemented through a finite element code which incorporates accurate crack tip singularity elements and which allows efficient fracture propagation through the mesh. The validity of the proposed method is studied in an experimental program using structural models in two types of rock. INTRODUCTION In 1913, Inglis (8) presented the first solution for the stress field corresponding to the problem shown in Figure 1. Since then, his solution to this problem has served as a well-spring for various theories of fracture and failure. However, it is now recognized that the problem of predicting e due to brittle fracture generated by elliptical flaws under n has never been solved. No theory has claimed to completely predict the behavior shown in Figure 2. Complete solution of this problem necessitates predictions of: Fracture initiation load Point of initiation on the flaw Incremental fracture loads and path of propagation Effects of interactions between flaws or between flaws and boundaries In this paper, classical and new concepts of fracture mechanics are combined into a proposed method for the study of compressive fracture-induced failure of rock structures. Both the finite element method and analytical techniques are used to implement the method on the structural configuration shown in Figures 1 and 2. The validity of the proposed method is studied in an experimental program using two types of rock. THE STABLE FRACTURE PROBLEM Many rock ( 3, 7) and fracture (4) mechanicians have documented experimentally the initially stable nature of fracture under compression. That is, fracture initiation and rupture are not synonymous under compression. Structural failure requires the propagation of fractures from initial flaws. Consequently, the fracture initiation load is only the first item of interest in a complete solution. For example, the sequence of photographs in Figure 2 depicts the increasing load necessary to propagate Figure 1: Biaxially Loaded Plate Containing a Central Elliptical Notch. ( Figure available in fullpaper) fractures under compression. Here, rupture was due to fracture/fracture and fracture/ boundary interaction. Failure occurred at a maximum load equal to 2.7 times the fracture initiation load. The basic, unanswered question arising from the behavior shown in Figure 2 is: A t each increment of load, which field variables govern the onset, direction, and length of the corresponding fracture increment? Theoretical concepts from classical rock mechanics and fracture mechanics will now be surveyed for potential contributions to the answer to this question.
- Well Completion > Hydraulic Fracturing (1.00)
- Well Drilling > Wellbore Design > Rock properties (0.35)
- Reservoir Description and Dynamics > Reservoir Characterization > Reservoir geomechanics (0.35)
INTRODUCTION Determination of the mechanical properties of New Mexico rock salt is motivated by the intent of the U.S. Energy Research and Development Administration (ERDA) to isolate existing and anticipated radioactive waste permanently from the environment. As one of various concepts under consideration, ERDA plans to construct a Waste Isolation Pilot Plant (WIPP) for transuranic and high level waste disposal experiments in the salt beds of the Delaware Basin in southeastern New Mexico. To support the design of this facility, a rock salt test program was initiated under the management of Sandia Laboratories. The purpose of this paper is to report the progress of this effort. The nature of the available material and the total test schedule for rock salt from southeastern New Mexico is outlined here. This discussion is followed by a description of new apparatus which were developed for this program. Finally, some preliminary results will be presented and discussed. Special emphasis was placed on the design and calibrations of equipment after consideration of the particular characteristics of rock salt (e.g. Schmidt, 1937; Corps of Engineers, 1963; LeComte, 1965; Serata et al., 1972; Dreyer, 1972; Heard, 1972) and the resulting experimental requirements. All of the work which is described in this paper was carried out at Sandia Laboratories and at the University of New Mexico. However, some of the experimental features incorporated here are based on research which was previously conducted at the University of Utah under the sponsorship of theOak Ridge National Laboratory and RE/SPEC, Inc. MATERIAL DESCRIPTION All rock salt examined was taken from core from a study area located east of Carlsbad, New Mexico, where drill holes, AEC 7 and AEC 8 had been drilled previously under a contract to the Oak Ridge National Laboratory. From this core, representative sections were taken from depths between 305 m (1000 ft) and 853 m (2800 ft) for further identification. Mineralogy Most samples are predominantly halite (NaC1) (average 83 modal percent, range 47-98%) with lesser amounts of anhydrite (CaS04) (average 4%, range 0- 154), polyhalite (K2MgCa2(SO4). 2H20) (average4 %,range 0-18%), and clay and silt (average 9%, range 0-44%). However, the modal data in parenthesis indicates that impurities may amount to as much as 44%. Halite is always present in the core with the exception of anhydrite layers (e.g. 7-2521 and 7-2522) and polyhalite seams (8-2366.1). Anhydrite and clay and silt are generally present but polyhalite is commonly absent. Trace amounts of magnesite, gypsum, other silicates, opaque oxides and rarely other sulfate and chloride constituents have also been observed.
- Materials (1.00)
- Energy > Oil & Gas > Upstream (1.00)
- Government > Regional Government > North America Government > United States Government (0.68)
- North America > United States > Texas > Permian Basin > Delaware Basin (0.99)
- North America > United States > New Mexico > Permian Basin > Delaware Basin (0.99)
- Reservoir Description and Dynamics > Reservoir Characterization (1.00)
- Well Drilling > Wellbore Design > Wellbore integrity (0.61)
- Health, Safety, Environment & Sustainability > Environment > Waste management (0.54)
ABSTRACT Comparison is made between two methods for predicting the consolidation settlement of a rigid, cylindrical, permeable punch pressed on the flat surface of a deep layer of wet clay substrate. The first method distributes the loading due to the punch as if it was in contact with an elastic medium, while the second method rigorously solves the contact problem between the punch and the consolidating medium. The predicted settlements agree favorably. INTRODUCTION The mathematical model for analyzing the consolidation behavior of wet clayey media under load, according to the Terzaghi-Biot theory (Terzaghi, 1925), (Biot, 1941a; 1955) led to solutions, too numerous to mention, which predict the settlement of clayey soils loaded at their surfaces by combinations of normal and tangential forces. Later on, contact problems, due to rigid stamps of varying geometry pressed on consolidating media, were solved and reported (Heinrich and Desoyer, 1961), (Agbezuge and Deresiewicz, 1974; 1975), (Chiarella and Booker, 1975). In the present paper, two methods are employed in predicting the settlement of a rigid, cylindrical, permeable punch pressed on the flat surface of a liquid-saturated porous elastic half-space. The first method considers the loading due to the punch to be distributed in an elastic manner, while the second method solves the contact problem between the two bodies. Reasonable agreement between the two solutions will indicate that for engineering applications, the first method is preferable because it is less complex. A REVIEW OF THE GOVERNING EQUATIONS The original equations given by the Terzaghi-Biot theory for the isotropic medium (Terzaghi, 1925), (Biot, 1941a) are expressed in terms of potential functions (McNamee and Gibson, 1960) by ; (Mathematical equation available in full paper) The general solution of (1) for a half-space loaded normally on its surface by a symmetric pressure p(r,t), distributed over the radial coordinate r, leads to an expression for the surface deflection (Agbezuge and Deresiewicz, 1974); (Mathematical equation available in full paper) and G denotes the shear modulus for the medium. In deriving (2), the following assumptions are made: The potential functions are bounded; the pore liquid diffuses freely at the loaded surface; p varies slowly with t; Poisson's ratio is zero; and compressive stresses are positive. THE PROBLEM AND ITS SOLUTION Let us consider the settlement of a rigid, cylindrical, permeable punch pressed on the flat surface of a deep layer of homogeneous clay (Fig. 1). The first method, which is approximate, will assume that the loading between the two bodies is distributed in elastic fashion (Timoshenko and Goodier, 1970) by; (Mathematical equation available in full paper) CONSOLIDATION SETTLEMENT OF A PERMEABLE PUNCH where N is the normal contact force, a the contact radius, and H(t) the Heaviside unit step function. Inserting (5) into (3) and the ensuing expression into (2), we obtain: (Mathematical equation available in full paper)
ABSTRACT ABSTRACT Fracture toughness, KIC is measured for single-edge-notch specimens of Indiana limestone as a, function of hydrostatic pressure. Result of 11 tests are reported for which specimens were precracked in fatigue, jacketed with urethane, and loaded to failure under superimposed confining pressure. A comparison of fracture toughness values for unconfined tests in two specimen configurations lends support to the concept that fracture toughness is a material constant for Indiana limestone. N Many failure criteria and theories such as the well-known Mohr criterion exist that predict failure conditions for rock. However, these theories often deal only with competent rock and do not deal directly with fracture processes. As a result, they cannot be expected to deal with questions of crack propagation such as (1) the length of major crack extension in processes such as hydraulic fracturing, (2) failure loads necessary to extend pre-existing fractures as might be encountered in repeated borehole explosive tests, and (3) conditions for crack bifurcation or branching necessary to create rubble in a rock fragmentation process. In the past 20 years, many investigations involving crack propagation in brittle metal alloys have employed the well-founded discipline of linear elastic fracture mechanics, LEFM, with great success. Although this theory is based. on linear elasticity and is directly related to the Griffith Theory (Griffith, 1924), plastic flow and other nonlinear behavior can occur on a small scale without affecting its predictive success. Purely brittle behavior is not required and only when the size of the zone of nonlinear behavior at a crack tip cannot be considered small when compared to the crack length does recourse to more exotic fracture theories such as the J integral (Rice, 1968) become necessary. LEFM is based on the stress intensity factor, K, which quantifies the intensity of the stress singularity at a crack tip. Fracture mechanics states that a. crack will advance when its stress intensity reaches a critical value, KIC, assuming that the crack tip is in a state of plane strain. This value of KIC , known as plane strain fracture toughness, has been shown to be a measurable material constant for a vast number of metal alloys, glasses, polymers, and even some organic materials such as wood, paper, and rubber. Fracture mechanics is related to Griffith theory (Griffith, 1924) which as modified by Orowan (Orowan, 1952) and restated by Irwin (Irwin, 1957) equates the critical rate of strain energy release during crack extension, GIc to twice the effective surface energy. (mathematical equation)(available in full paper) The effective surface energy is considered to include dissipative energy processes such as plastic flow and microcracking several rock mechanics investigations (Gilman,1960; Perkins and Krech, 1966; Friedman 1., 1972) have sought to measure yeff directly. Difficulties arise, even if plastic flow can be neglected, when attempts are made to estimate the total surface area created in crack propagation. On the other hand, the fracture mechanics approach has simply been to measure the value of the other side of the equation, the critical strain energy release rate, GIc by using a specimen with known crack geometry. (GIc is directly related to KIc, the critical stress intensity factor.)
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Geology > Rock Type > Sedimentary Rock > Carbonate Rock > Limestone (0.85)
- Well Completion > Hydraulic Fracturing (1.00)
- Reservoir Description and Dynamics > Reservoir Characterization > Reservoir geomechanics (0.48)
- Well Drilling > Wellbore Design > Rock properties (0.34)
INTRODUCTION An experimental program to study rock failure under the action of low velocity mechanical impact and locally applied static loads led to the need for a numerical technique to treat the time and spatial history of rock failure under such loading conditions. As such rock failure is concentrated under the point of impact or load application and yet is strongly influenced by the half-space nature of the rock body, a critical review of the various modeling techniques which might be utilized for such boundary conditions was carried out. In many other problems of rock deformation it would also be desirable to have a modeling technique which could treat locally nonelastic, rock failure in a half-space which was essentially subjected to elastic deformations. In addition to the application of an axisymmetric load on a half-space leading to failure (punch), examples of such problems are the application of a plane geometry load on a half space (wedge), the build-up of surface stresses leading to rock bursts during the removal of surface layers in a rock body with large residual strains, and the deformation along shear zones such as occurs with fault slippage and earthquake generation. Due to the nonlinear nature of rock failure and the complicated way in which deformation progresses in a locally loaded half space, it was concluded that non analytic techniques utilizing some type of zoning of the loaded half space would be required. Such zoning would allow the material included in each zone or element to be treated independently and arbitrarily in terms of permanent strain, failure criteria, residual strength during failure, dilatancy, etc. There are several techniques which were considered for possibly coupling the deformation and stresses in each discrete zone with the surrounding zones. Both conventional finite-element and finite-difference techniques could be used for problems of this type. For problems in which the deformations are not controlled by inertial effects, the finite-element technique would be, at first appearance, the most attractive of these two techniques. Modifications have been made, however, to the finite-difference technique utilized for dynamic problems such that static or kinematic problems can also be treated (Cassel and Hobbs, 1970; Andrews and Hancock, 1975). The biggest limitation of both the finite-element and finite-difference approaches derives from the fact that it is desirable to model the deformation of a crushed zone acting in a true half space. Both the finite-element and finite-difference techniques are limited in that all regions to be included in the analyses must be included in the zoning utilized. For a half-space problem this implies zoning to infinity which must somehow be compromised while still giving a reasonable representation of the physics involved.
- Geology > Geological Subdiscipline > Geomechanics (1.00)
- Geology > Structural Geology > Tectonics > Plate Tectonics > Earthquake (0.54)